Methods for leaching of ores

ABSTRACT

Disclosed and claimed are efficient methods for leaching minerals from ores using an acidic solution such as sulfuric acid. Additional factors which can improve mineral recovery include the use of an alkali metal halide, grinding the ore, addition of a carbon source, and/or adjustment of the temperature at which the process is carried out. Minerals such as titanium, iron, nickel, colbalt, silver and gold may be recovered by the methods of the present invention.

CROSS REFERENCE TO RELATED APPLICATION

[0001] This application is a continuation of co-pending application U.S.Ser. No. 09/507,547, filed Feb. 18, 2000. The subject application alsoclaims the benefit of the filing date of U.S. provisional applicationSer. No. 60/120,820 filed Feb. 19, 1999.

BACKGROUND OF THE INVENTION

[0002] Oxides of cobalt (Co), nickel (Ni), titanium (Ti), copper (Cu),molybdenum (Mo), lead (Pb), zinc (Zn), gold (Au), and silver (Ag) areimportant minerals. Various methods exist for recovering these compoundsfrom the ores where they are found. For example, autoclave methods areoften used to recover Co, Ni and Ti oxides. These methods are capitaland labor intensive. Mo oxide has been leached by hydrochloric acidmethods. Cyanide, thiosulfate, thiourea and halides are used in leachingAu and Ag metals and oxides. Cu, Zn and Pb can be leached with sulfuricacid.

[0003] Rutile (TiO₂) is a mineral used for many purposes. Amongst otheruses, it is a source of titanium metal and a paint pigment. Syntheticrutile is generally considered as any rutile created from anothermineral, usually ilmenite, that has at least 90% TiO₂. High purityrutile is 99.9%+ TiO₂. High purity rutile generally carries a commercialvalue premium.

[0004] Ilmenite (FeTiO₃) is most often converted to synthetic rutile byhigh temperature leaching with hydrochloric acid in an autoclave.Leaching temperatures are generally between 800 to 900° C. Ferricchloride is sometimes used in these autoclave leaches to increase thereaction rates at the lower temperatures.

[0005] Zoumei Jin et al (B. Mishra and G. J. Kiporous eds, In: TitaniumExtraction and Processing, The Mineral, Metals & Materials Society(1997) pg 122-128) reported that 4 to 6 Normal hydrochloric acid at 110to 140° C., will dissolve the iron (Fe) from ilmenite from the Sichuanprovince of China in 6 hours. They found that the reaction rate is 0.4order with respect to initial Fe⁺² concentration. They postulate asurface reaction control model with an apparent activation energy of56.97 kilojoules per mol.

[0006] Conventional autoclave technology is capital, maintenance andenergy intensive. The process disclosed in Zoumei Jin et al. processinvolves the use of large amounts of hydrochloric acid, which isexpensive, difficult to handle and requires special stainless steelequipment. There is a clear need for more efficient processes forleaching of ores to obtain valuable minerals.

[0007] Cyanide is the most commonly used leachant for gold. Twomolecules of cyanide complex with every molecule of gold. Copper alsocomplexes with cyanide, but it takes 4 molecules of cyanide for everycopper molecule. Copper is often present in copper gold ores in the onetenth to one percent range. Gold in these ores is in the one to 10 partsper million range. The copper consumes so much cyanide it needs to berecovered by hydrogen cyanide distillation, an expensive and dangerousoperation. Systems have been proposed where sulfuric acid is used toleach the copper first. Then, the heap or batch is neutralized. Cyanidecan then be used to leach the gold. The problem, of course, is theexpense of neutralization. In heap operations, the additional worry ofincomplete neutralization is present.

[0008] In other ores, the gangue, or unwanted material, can be an acidconsumer. Copper oxide in limestone rich rock is an example.

BRIEF SUMMARY OF THE INVENTION

[0009] The subject invention pertains to novel and highly efficientmethods for leaching valuable minerals, such as cobalt (Co), nickel(Ni), titanium (Ti), copper (Cu), molybdenum (Mo), lead (Pb), zinc (Zn),gold (Au), and silver (Ag) from ores.

[0010] One aspect of the present invention concerns methods forrecovering titanium from ores. One embodiment of the subject method usesan acidic solution, such as sulfuric acid, to leach titanium oxides froma mineral feed. Additional modifications and/or steps, including, forexample, grinding of the ore, addition of an alkali metal halide,addition of a carbon source, and adjustment of pressure and/ortemperature, can be incorporated in the process. In a preferredembodiment, a mineral feed is contacted with an acid and an alkali metalhalide to leach titanium oxides from the feed. High purity titaniumdioxide having a commercial premium over synthetic rutile can beproduced using the methods of the subject invention.

[0011] Another aspect of the present invention concerns methods forrecovering transition metals other than titanium from ores. In oneembodiment, the present invention provides a method for recovery ofnickel and cobalt from a mineral feed by leaching the feed with anacidic solution. In an exemplified embodiment, a mixture of sulfuricacid and an alkali metal halide are used to leach out cobalt and nickelfrom a laterite ore. The subject methods can also be used to recovercobalt, nickel, copper, etc. by leaching these elements from scrapmetal.

[0012] The subject invention also concerns methods for recoveringmultiple metals or metal oxides in separate solutions. In oneembodiment, ore is contacted with an acid solution, such as sulfuricacid. Solid residue is then collected and contacted with an alkali metalhalide and acid solution. In an exemplified embodiment, the subjectmethod is used to recover copper separately from gold and silver. Thecopper is recovered primarily in the first acid solution, while the goldand silver are recovered in the alkali metal halide and acid solution.

BRIEF DESCRIPTION OF THE DRAWINGS

[0013]FIG. 1 shows the kinetics of leaching titanium and iron fromilmenite.

[0014]FIG. 2 shows the results of four consecutive one-hour leaches oftitanium and iron from ilmenite.

[0015]FIG. 3 shows pulp density relationships in the leaching oftitanium and iron from ilmenite.

[0016]FIG. 4 shows the results of experiments evaluating the effect ofan alkali metal halide (NaCl) on the sulfuric acid leaching process.

[0017]FIG. 5 shows the results of experiments evaluating the effects ofgrinding the ore on recovery rates.

[0018]FIG. 6 shows the results of experiments evaluating the effect ofadding a carbon source during the sulfuric acid leaching process.

[0019]FIG. 7 shows the results of experiments evaluating the effect ofan alkali metal halide on the sulfuric acid leaching process of leachingnickel from an initial laterite feed (Laterite-1).

[0020]FIG. 8 shows the results of experiments evaluating the effect ofan alkali metal halide on the sulfuric acid leaching process of leachingcobalt from an initial laterite feed (Laterite- 1).

[0021]FIG. 9 shows the results of experiments evaluating the effect ofan alkali metal halide on the sulfuric acid leaching process of leachingnickel from a second laterite feed (Laterite-2).

[0022]FIG. 10 shows the results of experiments evaluating the effects ofan alkali metal halide on the sulfuric acid leaching process of leachingcobalt from a second laterite feed (Laterite-2).

DETAILED DISCLOSURE OF THE INVENTION

[0023] The subject invention provides novel materials and methods usefulfor the recovery of minerals from ores. An important component of theleaching processes of the subject invention is the use of an acidicsolution. In one embodiment, the acid is sulfuric acid. Sulfuric acidused in the leaching procedures can be at a concentration ranging fromabout 20 grams per liter to about 500 grams per liter. In a preferredembodiment, the concentration of sulfuric acid ranges from about 150grams per liter to about 250 grams per liter. Preferably, theconcentration of sulfuric acid is approximately 200 grams per liter.

[0024] In addition to using sulfuric acid solutions in the leachingprocesses of the subject invention, particularly preferred embodimentsof the subject invention utilize additional factors including, forexample, the use of an alkali metal halide, grinding the ore, additionof a carbon source, and/or adjustment of the temperature at which theprocess is carried out.

[0025] In accordance with the subject invention, the efficiency of theleaching process can be improved by grinding the ore prior to treatment.In a preferred embodiment, the ore is ground so that it can pass througha 200 mesh sieve.

[0026] In a further embodiment, an alkali metal salt can be added to theleach solution to improve recovery. The alkali metal salt can be forexample, an alkali metal halide, alkali metal nitrite, alkali metalnitrate, alkali metal sulfite or alkali metal thionite. The metal halidecan be, for example, NaCl, KCl, NaBr or KBr, or mixtures of one or moreof these. The metal sulfites can be, for example, sodium sulfite, sodiummetabisulfite, sodium bisulfite, sodium dithionite, or other alkalimetal or ammonium sulfite, metabisulfite, bisulfite or dithionite. Theordinarily skilled artisan, having the benefit of the teachingsdisclosed herein, can readily determine those alkali metal salts thatcan be used in conjunction with the particular acid solution used in thesolubilization step of the process.

[0027] A further embodiment of the subject invention involves the use ofa carbon source to improve recovery. The carbon source can be, forexample, graphite or activated carbon. The source of this material canbe, for example, from coconut shell or wood.

[0028] The present invention accordingly provides in one embodiment amethod for recovering titanium oxide(s) from a titanium andiron-containing mineral feed, the method including the steps of:

[0029] (a) solubilizing titanium and iron by leaching the feed with anacidic solution in the presence of an alkali metal halide;

[0030] (b) selectively precipitating titanium oxide(s), and

[0031] (c) recovering titanium oxide(s).

[0032] Typically, the titanium oxide(s) may be titanium dioxide.

[0033] The titanium-containing mineral feed is typically post heavymineral concentration products. The feed will include titaniummineralization. Typical examples of this titanium mineralization areilmenite (FeTiO₃), leucoxene, perovskite (CaTiO₃) and titano magnetite.

[0034] The feed may in an alternative embodiment comprise a bulkilmenite concentrate. Other titanium-containing mineral feed materialsare contemplated within the scope of the invention.

[0035] The present invention provides in another separate embodiment amethod for recovering synthetic rutile (TiO₂), from a mineral feedcomprising ilmenite (FeTiO₃), the method including the steps of:

[0036] (a) solubilizing titanium and iron by leaching the ilmenite withan acidic solution in the presence of an alkali metal halide;

[0037] (b) selectively precipitating titanium oxide(s), and

[0038] (c) recovering titanium oxide as TiO₂.

[0039] In step (a) of the method, the acidic solution preferablyincludes sulfuric acid. The sulfuric acid used in the leaching step istypically at a concentration in the range of from about 20 grams perliter to about 500 grams per liter. In a preferred embodiment, theconcentration of sulfuric acid is in the range of from about 150 gramsper liter to about 250 grams per liter. Most preferably theconcentration of sulfuric acid is about 200 grams per liter. Other acidscontemplated for use in step (a) of the present invention include, butare not limited to, a halide acid such as hydrochloric acid orhydrobromic acid. The typical concentration of halide acid used is inthe range of from about 150 to about 350 grams per liter.

[0040] Step (a) is typically carried out in the presence of an alkalimetal halide at a ratio of alkali metal halide to ilmenite in the feedin the range of from about 1:1 to 2:1. Preferably, the ratio of alkalimetal halide is from about 1:1 to 1.5:1. More preferably, the ratio isabout 1.2:1. Suitable alkali metal halides include, but are not limitedto, NaCl, KCl or KBr or mixtures of one or more of these.

[0041] In the methods of the present invention, the alkali metal halidecan be added directly to the leach solution. Alternatively, the alkalimetal halide can be combined with the feed prior to introduction of theleaching solution. In this case, the feed may be subjected to a boildownpretreatment (i.e., by boiling down to approximate dryness) in thepresence of the alkali metal halide whereby the feed (e.g., ilmenitesurfaces) are coated with alkali metal halide prior to leaching.Optionally, a combination of the foregoing, i.e., direct addition ofalkali metal halide to the feed and combination of alkali metal halidewith the feed prior to leaching, can be used in the subject methods.Thus, for example, a proportion of the alkali metal halide is combinedwith the feed prior to step (a) and a proportion of the alkali metalhalide is added directly to the leach solution. Typically, steps (a) and(b) may be conducted simultaneously or separately once solubilizationcommences. It is particularly preferred to concurrently remove some ofthe pregnant solution from the leach residue to permit precipitation totake place away from the leach residue. In this way, the precipitate maybe restricted from coating the leach residue which could potentiallydecrease the efficiency of the process.

[0042] In one embodiment, the precipitation step (b) can be regulated byadjustment of temperature and/or pH of the solution. Typically, step (a)is carried out at a temperature in the range of from about 80° C. toabout 120° C. and, preferably, is in the range of from about 90° C. toabout 110° C. In a preferred embodiment, the operating temperature forstep (a) is about 100° C.

[0043] In one embodiment, the leach solution in step (a) has a solidscontent of up to about 60% by weight. Preferably, the leach solution hasa solids content of from about 10% to about 40%.

[0044] To facilitate more rapid leaching, the feed may be ground intofiner particles. In a preferred embodiment, the feed may be subjected tofine grinding. Preferably, the majority of particles in the feed arecapable of passing through a 75 micron sieve after grinding.

[0045] Optionally, a source of carbon may be provided in the subjectmethod. The carbon may be in the form of any commercially availablecarbon source including, for example, activated carbon, coal, coke,charcoal or graphite. A preferred source of carbon is activated carbonderived from coconut shell. The ratio of carbon to feed (e.g., ilmenite)is typically between about 0.01:1 to 1:1.

[0046] Methods according to the present invention may be carried out ator above atmospheric pressure. When elevated pressures are used, thetypical elevated pressures and temperatures at which the present methodsmay be performed are in the range of from about 1 bar to about 30 bar.Preferably, pressures are in the range of from about 1 bar to about 5bar. Temperatures used in the subject methods range from about 100° C.to about 235° C. Preferably temperatures range from about 100° C. toabout 150° C.

[0047] The leach residue produced from step (a) can be subjected tofurther leaching to solubilize undissolved iron and/or titanium in theresidue. The further leaching can be performed using fresh acidicsolution. In an alternative embodiment, spent leach liquor or acombination of fresh acidic solution and spent leach liquor, can beused.

[0048] In another embodiment, step (a) of the subject method can beperformed in the presence of ferrous and/or ferric ions to promotedissolution of the iron mineralization. Ferrous ions will generally bepresent in recirculated process plant solutions.

[0049] If desired, iron may be removed from the leachant solution usingstandard techniques, such as precipitation. The purpose is to removesoluble iron from any process solutions. Solvent extraction, ionexchange, reverse osmosis or other techniques can also be used to removesoluble iron.

[0050] The leach time for this embodiment is generally relatively long,and typically is in the range of from about 50 to about 120 hours.Preferably, leach time is from about 60 to about 100 hours. However, theoperating conditions are much milder than conventional autoclavetechniques, leading to large capital and operating cost advantages.Sulfuric acid and alkali metal halides are easier to handle than thehydrochloric acid used in the Zoumei Jin et al. process referred toabove.

[0051] The present invention provides in another separate embodiment, amethod for recovering titanium from a titanium and iron-containingmineral feed, the method including the steps of:

[0052] (a) solubilizing titanium and iron by leaching the feed with anacidic solution in the presence of an alkali metal halide and a sourceof activated carbon;

[0053] (b) selectively precipitating titanium oxide(s), and

[0054] (c) recovering titanium oxide(s) from the leach residue.

[0055] The present invention provides in another separate embodiment amethod for recovering titanium from a mineral feed comprising ilmenite(FeTiO₃), the method including the steps of:

[0056] (a) solubilizing titanium and iron in the ilmenite by leachingthe ilmenite with an acidic solution in the presence of an alkali metalhalide and a source of activated carbon;

[0057] (b) selectively precipitating titanium oxide(s), and

[0058] (c) recovering titanium oxide from the leach residue as TiO₂.

[0059] The present invention provides in another separate embodiment amethod for recovering titanium oxide(s) from a mineral feed comprisingilmenite (FeTiO₃), the method including the steps of:

[0060] (a) leaching the ilmenite with an acidic solution at atemperature in the range of from about 80 to 120° C. in the presence ofan alkali metal halide for a predetermined time, the leach solutioncontaining up to about 60% by weight solids to produce a leachantsolution containing iron and titanium ions;

[0061] (b) separating the iron from the titanium in the leachantsolution; and

[0062] (c) recovering the separated titanium as TiO₂.

[0063] As mentioned above, maintaining the titanium in solution ratherthan allowing it to report to the residue as a precipitate has beenobserved to further enhance the likelihood of the titanium beingrecovered as a pure product. Where most of the titanium reports to theresidue, other materials that may be found in proximity with theilmenite mineral including chromite, lime, magnesia, silica orsilicates, manganese, alumina, vanadium, phosphate and zirconium willalso tend to remain in the residue along with undissolved iron. Thepresence of such materials is likely to dilute the purity of titaniumrecoverable from the residue.

[0064] Depending on the metals content of the leach solution, a typicalreaction time for step (a) of this embodiment is up to about an hour.Preferably, the reaction time of step (a) is up to about half an hour.More preferably, the reaction time is in the range of from about 5 toabout 15 minutes. It has been observed that titanium solubility reachesa peak during reaction times of approximately that length.

[0065] Optionally, step (a) above may be repeated to solubilizeunleached titanium in the residue obtained following step (a) in orderto obtain cumulative maximum solubility of titanium. Fresh acidicsolution and alkali metal halide can be used when step (a) is repeated.Step (a) may in one embodiment comprise a type of countercurrent leachcircuit.

[0066] The acidic solution in this embodiment can be supplemented withhydrochloric acid in one or more steps of a repeated leach sequence toassist in enhancing the titanium solubility profile.

[0067] In another separate embodiment the present invention provides amethod for recovering titanium from a titanium and iron-containingmineral feed, the method including the steps of:

[0068] (a) contacting the feed material with an halide acid solution oran acid—alkali halide solution for a period of time sufficient tosolubilize the titanium but insufficient to allow the titanium toappreciably precipitate;

[0069] (b) selectively precipitating titanium oxide(s); and

[0070] (c) recovering titanium oxide(s).

[0071] The halide acid used in step (a) can be, for example,hydrochloric acid or hydrobromic acid. The concentration of halide acidused can be in the range of from about 150 to about 350 grams per literacid.

[0072] Any precipitated titanium reporting to the leach residue of thisembodiment may be recovered in subsequent leaching operations.

[0073] The present invention provides in another separate embodiment amethod for recovering titanium from a feed comprising finely groundilmenite (FeTiO₃), the method including the steps of:

[0074] (a) leaching the ilmenite with an acidic solution containingsulfuric acid at a temperature of about 100° C. in the presence of analkali metal halide selected from the group consisting of NaCl, KCl andKBr and in the presence of a source of activated carbon for up to abouthalf an hour to produce a leachant solution containing iron and titaniumions, the ratio of alkali metal halide to ilmenite in the feed beingabout 1.2:1; and the ratio of activated carbon to ilmenite in the feedbeing about 0.01:1, the solids content of the leach solution being up toabout 60% by weight;

[0075] (b) repeating step (a);

[0076] (c) separating at least some of the pregnant solution from theleach residue;

[0077] (d) selectively precipitating the titanium oxide(s) from thepregnant solution; and

[0078] (e) recovering the titanium oxide as TiO₂.

[0079] In a particularly preferred embodiment, the present inventionprovides multistage leaching of iron and titanium from an iron andtitanium-bearing mineral feed, the method comprising the followingsteps:

[0080] (a) contacting the feed material with an acid—alkali halidesolution for a period of time sufficient to solubilize the titanium butnot so long as to allow the titanium to appreciably precipitate;

[0081] (b) separating the pulp from the leach liquor;

[0082] (c) contacting the pulp with fresh leach liquor and repeatingsteps (a) and (b) until all economically feasible titanium is leached;and

[0083] (d) selectively recovering the titanium and iron in separatestages from the leach solutions by precipitation, solvent extraction orother means.

[0084] The conditions of step (a) can involve percent solids on aweight/weight basis of between about 1 percent and about 60 percent. Thetypical percent solids are in the range of from about 10% to 40%. Thesolids may be ground to fine size to facilitate leaching, typically sothat the feed passes a 73 micron sieve. The acid used is most typicallysulfuric acid. The acid concentration can range from about 20 to about300 grams per liter acid. Most typically the acid concentration rangesfrom about 150 to 230 gram per liter.

[0085] The alkali halide can be any alkali halide. Preferably, thealkali halide is NaCl, KCl, NaBr, or KBr. The concentration of alkalihalide can range from about 50 grams per liter to about 400 grams perliter. Preferably, the alkali halide concentration is about 100 to about200 grams per liter.

[0086] The leaching is most typically carried out at about roompressure. The temperature can be between about 40° C. and about 110° C.at room pressure. Preferably, leaching temperature is between about 90°C. and about 105° C. Leaching at room pressure will typically beperformed in a leach vessel with a condenser to limit the loss of halideacid generated in the leach solution. The titanium is allowed to reach aconcentration as high as possible before it begins to re-precipitateonto the leach feed material. This is typically slightly over four (4)grams of titanium per liter of solution. The leach time to accomplishthis solubilization will depend on the various aforementioned parametersbut will usually range from about 10 minutes to 1 hour.

[0087] The method of solid—liquid separation in step (b) can be anymethod that makes a good separation of the solids from the leach liquorin a relatively short time. These include methods such as cyclones,filters, centrifuges, magnetic separators, and settlers. The list is notmeant to exclude any unnamed method.

[0088] The fresh leach liquor in step (c) can be leach liquor from whichthe titanium content has been reduced or eliminated. The iron content ofliquor should be controlled so that no precipitation of an iron compoundoccurs during the leaching.

[0089] The titanium can be totally or partially removed from the leachliquor in step (d) by the method that makes the most economic sense forany given plant. The methodology available includes, but is not limitedto, precipitation by seeding or pH adjustment, crystallization, solventextraction, and ion exchange.

[0090] The iron can be removed in a similar fashion in a step before orafter the titanium recovery. Titanium and iron are recovered as separateproducts, in separate stages. The titanium would be recovered as atitanium salt, most typically TiO₂. The iron would most typically berecovered as an iron salt such as ferrous chloride or sulphate.

[0091] In addition to titanium and iron leaching, the present inventionalso concerns methods for the recovery of other minerals, such asnickel, cobalt, copper, molybdenum, lead, zinc, gold or silver from ore,soil, concentrate, slag or residue. In one embodiment, a method isprovided for the dissolution of nickel and cobalt from a nickel, cobaltand iron-containing mineral feed, the method comprising solubilizing thenickel, cobalt and iron in the feed by leaching the feed with an acidicsolution. In a further embodiment, an alkali metal salt can be added tothe leach solution to improve recovery. The alkali metal salt can be forexample, an alkali metal halide, alkali metal nitrite, alkali metalnitrate, alkali metal sulfite or alkali metal thionite. The metal halidecan be, for example, NaCl, KCl, NaBr or KBr, or mixtures of one or moreof these. The metal sulfites can be, for example, sodium sulfite, sodiummetabisulfite, sodium bisulfite, sodium dithionite, or other alkalimetal or ammonium sulfite, metabisulfite, bisulfite or dithionite. Theordinarily skilled artisan, having the benefit of the teachingsdisclosed herein, can readily determine those alkali metal salts thatcan be used in conjunction with the particular acid solution used in thesolubilization step of the process. In another embodiment, the method ofthe invention can be conducted at above ambient temperatures and at orabove atmospheric pressures prior to metal extraction by precipitation,solvent extraction or other means.

[0092] Where the metals of interest are nickel and cobalt, the nickeland cobalt-containing mineral feed is typically post beneficiation bycomminution and thickening products. A typical example of nickel andcobalt mineralization is a laterite ore. Alternatively, the feed maycomprise a bulk laterite concentrate.

[0093] One embodiment of the present method provides for recoveringnickel and cobalt from a mineral feed comprising laterite, the methodincluding the step of solubilizing nickel and cobalt and iron in thelaterite by leaching the laterite with an acidic solution in thepresence of an alkali metal halide at a temperature not exceeding about150° C. at normal pressures prior to nickel and cobalt extraction byestablished precipitation, solvent extraction or other means.

[0094] Preferably, the acidic solution contains sulfuric acid. Thesulfuric acid used in the leaching step is typically at a concentrationin the range of from about 20 grams per liter to about 500 grams perliter. In a preferred embodiment, the concentration of sulfuric acid isin the range of from about 150 grams per liter to about 250 grams perliter. Preferably, the concentration of sulfuric acid is about 200 gramsper liter. Other acids contemplated for use in the present inventioninclude halide acids, for example, hydrochloric acid or hydrobromicacid. The typical concentration of halide acid used is in the range fromabout 50 to about 350 grams per liter acid.

[0095] The process is typically carried out in the presence of an alkalimetal halide at a ratio of alkali metal halide to laterite in the feedin the range of from about 0.05:1 to about 4:1. Preferably, the ratio isabout 0.1:1, and most preferably about 0.2:1.

[0096] In any of the described embodiments of the invention, includingthose methods directed towards leaching of titanium and non-titaniumtransition elements from a mineral feed, the alkali metal salt may beadded directly to the leach solution. Alternatively, the alkali metalsalt is combined with the feed prior to introduction of the leachingsolution. In this case, the feed may be subjected to a boildownpre-treatment (i.e., by boiling down to approximate dryness) in thepresence of the alkali metal salt whereby the feed (e.g., laterite)surfaces are coated with alkali metal salt prior to leaching. In anotheralternative embodiment the solution of alkali salt may be sprayed on aheap of lateritic ore and allowed to evaporate. Further a combination ofthe foregoing may be adopted. Namely, a proportion of the alkali metalsalt is combined with the feed prior to solubilization and a proportionof the alkali metal salt is added directly to the leach solution. It isparticularly preferred to concurrently remove some of the pregnantsolution from the leach residue to permit separation of the nickel andcobalt to take place away from the leach residue.

[0097] Typically, the process is carried out at a temperature in therange of from about 80° C. to about 120° C. Preferably, the temperatureis in the range of from about 90° C. to about 110° C. A typicaloperating temperature for the process is about 100° C.

[0098] The leach solution in the subject process preferably has a solidscontent of up to about 60% by weight. Preferably, the leach solution hasa solids content of from about 10 to 40%.

[0099] To facilitate rapid leaching, the feed can be ground into smallerparticles. It is preferred that the feed be subjected to fine grinding.Preferably, the majority of particles in the feed are capable of passingthrough 75 micron sieve. Typically, at least 75% of the particles in thefeed are of a size that can pass through 75 micron sieve openings.

[0100] Methods according to the present invention may be carried out ator above atmospheric pressure. When elevated pressures are used, thetypical elevated pressures and temperatures at which methods accordingto the invention may be performed are in the range of from about 1 barto about 30 bar. Preferably, pressures are in the range of from about 1bar to about 5 bar and temperatures range from about 100° C. to about235° C. Preferably in the range of from about 100° C. to about 150° C.The methods described in the embodiments of the present invention do notconflict with known autoclave technology as the present inventioninvolves the use of alkali metal halides in combination with sulfuricacid whereas known autoclave technology utilizes pure acid or ammoniacalsolutions to leach the nickel and cobalt from lateritic feed ores.

[0101] The leach residue produced by the present process may besubjected to further leaching to solubilize undissolved iron and/ornickel and cobalt in the residue. The further leaching can be performedusing fresh acidic solution. In an alternative embodiment, spent leachliquor, or a combination of fresh acidic solution and spent leachliquor, may be used in the process.

[0102] Additionally, the process may be performed in the presence offerrous and/or ferric ions to promote dissolution of the ironmineralization. Ferrous ions will generally be present in recirculatedprocess plant solutions.

[0103] Depending on the metals content of the leach solution, a typicalreaction time for the process of this embodiment is up to about sixhours. Preferably, the reaction time is up to about two hours. Morepreferably, the reaction time is in the range of from about 15 minutesto about 3 hours. It has been observed that nickel and cobalt solubilityreaches a peak during reaction times of approximately that length. Aperson of ordinary skill in the art can vary leach time so as to leachless of an unwanted species such as manganese or iron at the expense ofsome cobalt and nickel recovery.

[0104] The process above may be repeated to solubilize unleached nickelor cobalt in the residue in order to obtain cumulative maximumsolubility of nickel and cobalt. Fresh acidic solution and alkali metalhalide may be used when the process is repeated. The process may in oneembodiment comprise a type of countercurrent leach circuit.

[0105] The acidic solution may in this embodiment be supplemented withhydrochloric acid in one or more steps of a repeated leach sequence toassist in enhancing the nickel or cobalt solubility profile.

[0106] In another embodiment of the present invention, a metal halidesalt may be used either to precondition an aqueous slurry or it may besprayed onto the feed material and allowed to evaporate prior tocontacting with sulfuric acid.

[0107] Upon contact with sulfuric acid the resultant slurry is permittedto leach for a short time (typically less than about fifteen minutes)but most preferably about five minutes or less. The liquid is thenseparated and sent for cobalt recovery. This flash leaching processutilizes the selective nature of the leach to achieve a cobalt richsolution containing only minor amounts of nickel, manganese, iron, etc.

[0108] The residue from the flash leach is subsequently leached with themetal halide sulfuric acid mixture for longer periods of time tosolubilize the nickel and any remaining cobalt.

[0109] In another embodiment super alloy scrap and other recycled metalalloys may be leached by treating with a halide salt of the metal andsulfuric acid. The concentrations of the metal halide salt and thesulfuric acid will be dependent upon the specific scrap mixture. Thisembodiment can be utilized to selectively leach specific metals or toplace all the metals into solution. This embodiment may also be used tosolubilize radio-nucleosides of such metal as nickel from a radiatedscrap. Oxygen or other oxidizing gasses such as chlorine can be added tothe system to oxidize the metal.

[0110] For some oxide ores containing minerals that contain multivalenttransition metals such as Co and Mn in an high oxidation state species,the alkali metal halide may be substituted with a sulfur-based reducingchemical. For example, sodium sulfite, sodium metabisulfite, sodiumbisulfite, sodium dithionite, or other alkali metal or ammonium sulfite,metabisulfite, bisulfite or dithionite can be used in place of thealkali metal halide. These sulfur based reducing chemicals willfacilitate the reduction of the transition metal, opening the ore up toattack by the sulfuric acid. The metal of economic interest need notalways be the one reduced. Alkali metal nitrates or nitrites may be usedwith sulfuric acid to leach most metals. These techniques may be used toleach metals from sulfide minerals or from scrap, residue, slags,concentrates, or soils.

[0111] In another embodiment the process utilizing a metal halide saltand sulfuric acid may be used, with minor modifications, in currentlyexisting counter current decantation (CCD) circuits. Such an embodimentwould utilize fresh feed material to achieve neutralization to a pHadequate to retain iron in solution. After a liquid-solid separation hasbeen effected, the resultant leach liquor may be further neutralized toprecipitate iron as a hydroxide in the presence of a binding material.The iron precipitate may then be partially dried and pelletised toproduce pig iron feed stocks.

[0112] The method of solid-liquid separation can be any method thatproduces a good separation of the solids from the leach liquor in arelatively short time. These include, but are not limited to, methodssuch as cyclones, filters, centrifuges, magnetic separators, andsettlers.

[0113] The nickel or cobalt can be totally or partially removed from theleach liquor by the method that makes the most economic sense for anygiven plant. The methodology available includes, but is not limited to,precipitation by seeding or pH adjustment, crystallization, solventextraction, and ion exchange.

[0114] The subject invention also concerns methods for recoveringmultiple metal or metal oxides in separate solutions. Mineral species ofeconomic value are often associated with species that consume thechemical reagents that are used to leach them. Sometimes even though theconsuming species is of economic value, the overall leach becomesuneconomic. The most common example of this is copper-gold ores.

[0115] Following are examples which illustrate procedures for practicingthe invention. These examples should not be construed as limiting. Allpercentages are by weight and all solvent mixture proportions are byvolume unless otherwise noted.

EXAMPLE 1 Leaching of Titanium and Iron from Ilmenite

[0116] Kinetics experiments on the leaching of titanium and iron fromilmenite shows that both are leached early on and that the titanium thenprecipitates and slows the iron leaching. An experiment with 100 gramsof ilmenite, ground to −200 mesh was conducted. The tests were conductedwith 1000 grams 200 gram per liter (g/l) sulfuric acid with 120 g/l NaClsolution. 100 grams activated carbon were added and the solution heatedto 100° C. The Fe and Ti concentrations were monitored during the courseof the 96 hour leach. The results are presented in FIG. 1. The resultspresent a mechanism of initial Ti leaching into the liquor withsubsequent hydroxylation and subsequent precipitation. While this occursit slows leaching of the iron. The Ti appears to be leached within onehour.

[0117] In a separate but analogous experiment a 100 gram quantity ofilmenite with a head assay of 34.0% Fe and 27.0% Ti, and particle sizesuch that 100% of the particles pass through a 75 micron screen, wasleached for 72 hours at 100° C. in 1 liter of 200 gram per literH₂SO₄—120 gram per liter alkali metal halide solution. A 100 gramquantity of activated carbon was also present in the leach solution. Theleach liquor was monitored periodically for Ti and Fe content. Theresults of the experiment are shown in Table 1. Titanium is dissolvedthen observed to subsequently precipitate. The final assay of the 57.4gram residue showed that it contained only 0.67% Fe and 46.6% Ti. Thus,98.9% of the iron had been extracted into the solution while 99.7% ofthe titanium remained in the residue. The experiment indicates that dueto the initial solubilization of Ti, both Ti and Fe can ideally beextracted from ilmenite by repeated short duration leaches. TABLE 1Concentration of Titanium in Residue by Dissolution of Iron fromIlmenite Fe Ti Time mg/l Cumulative Cumulative Hours Fe Ti Volume litersgm Extraction gm Extraction  1 5600 3720 0.720 4.03 11.9 268 9.9  2 56503700 0.720 4.07 12.3 266 10.0  4 5950 3810 0.720 4.28 13.1 274 10.6  66010 3880 0.720 4.33 13.4 279 10.9 12 6220 3830 0.720 4.48 14.0 276 10.924 16900 1410 0.720 12.17 37.1 1.02 4.5 48 35000 212 0.720 25.20 76.50.15 1.3 72 38200 121 0.720 27.50 84.4 0.09 1.1 96 37600 90 0.720 27.0798.9 0.06 1.0 Wash 1 5000 20 1.000 5.00 14.7 0.02 0.1 Wash 2 375 1 0.9900.37 1.1 0.00 0.0

EXAMPLE 2 Consecutive One-Hour Leaching of Titanium and Iron fromIlmenite

[0118] Using the data in Example 1 allowed the development of a newleach procedure for ilmenite. The procedure comprises leaching ilmenitefor one hour or less and then contacting it with fresh leach solution.In this manner both the iron and titanium is leached together. This wastested using the same conditions as in the 96-hour test. The results offour (4) consecutive one-hour leaches on the same ore sample are shownin FIG. 2. As can be seen, approximately the same amount of iron andilmenite was leached in each step. The ordinarily skilled artisan,having the benefit of the teachings described herein, can determine theproper reagent concentration, temperature, particle size of the ore,whether to include carbon and its form (e.g., activated carbon orgraphite), or atmospheric pressure (typically <3 atmospheres) that isoptimum for a particular ore. The technique of separating the Ti as TiO₂with short leach times followed by precipitation of TiO₂ is alsoapplicable to other leach systems such as the hydrochloric acid leachsystem.

[0119] The following two experiments further demonstrate methodology forleaching both the titanium and iron from ilmenite in a multistagefashion:

[0120] Experiment A comprises a leach solution of 60 grams alkali metalhalide, 100 grams H₂,SO₄, and 350 grams of H₂O heated to 100° C. inErlenmeyer flasks on a stirring hotplate to which is added 50 grams ofminus 75 microns particle size ilmenite resulting in a 9% pulp density.

[0121] Experiment B comprises a leach solution of 60 grams alkali metalhalide, 100 grams H₂SO₄, and 350 grams of H₂O heated to 100° C. inErlenmeyer flasks on a stirring hotplate to which is added 100 grams ofminus 75 microns particle size ilmenite resulting in a 16% pulp density.

[0122] The ilmenite had an assay head of 30% titanium and 34% iron.

[0123] The following procedure steps are applied separately toExperiment A and Experiment B:

[0124] Step 1. A condenser is placed on the Erlenmeyer containing theslurry comprising the prescribed solution and ilmenite feed;

[0125] Step 2. The slurry is stirred vigorously with a magnetic stirrerfor 30 minutes with the temperature maintained at 100° C.;

[0126] Step 3. The Erlenmeyer and contents are cooled for a couple ofminutes in a room temperature water bath;

[0127] Step 4. The Erlenmeyer solution is decanted into a centrifugerube and centrifuged at 4,000 rpm for 5 minutes;

[0128] Step 5. The liquor in the centrifuge tube is decanted andseparated from the solids into a sample bottle, volume and weightdetermined and retained for further test work including analysis;

[0129] Step 6. The remaining solids in the centrifuge tube are weighedand then washed, with 510 grams of fresh leach solution, back into theresidue remaining in the Erlenmeyer after Step 4;

[0130] Step 7. The reconstituted slurry is stirred and the slurrytemperature increased to 100° C.;

[0131] Step 8. The procedure is continued by repeating Steps 1 through 7inclusive, a total of seven times and thus equating to a total leachduration of 4 hours;

[0132] Step 9. The post centrifuging liquors collected at eachrepetition of Step 5 are individually subsampled and analysed fortitanium and iron;

[0133] Step 10. Calculations are conducted to determine titanium andiron contents of both solids and liquors and comparisons made with therespective elemental assay values of the ilmenite ore feed;

[0134] Step 11. The individual liquors remaining after the subsamplingconducted in Step 9 are combined in a flask and subsampled and analysedfor titanium and iron;

[0135] Step 12. The titanium can be totally or partially removed fromthe leach liquor by the method that makes the most economic sense forany given plant. The methodologies available include, but are notlimited to, precipitation by seeding or pH adjustment, crystallization,solvent extraction, and ion exchange.

[0136] The results of the experiments are shown in Table 2 and FIG. 3.For both levels of percent solids the trend is for roughly a constantamount of titanium to be extracted at each step. TABLE 2 Pulp DensityRelationships on the Leaching of Titanium and Iron from Ilmenite 9%Solids on a weight/weight basis g/l Fe Ti Time Volume CumulativeCumulative Leach Hours Fe Ti Liters gm Extraction gm Extraction  1 0.56.32 3.96 0.400 2.53 14.8% 1.58 10.6  2 0.5 3.97 3.32 0.423 1.68 24.71.40 19.9  3 0.5 3.14 2.88 0.412 1.29 32.3 1.19 27.8  4 0.5 2.15 2.180.410 0.88 37.4 0.89 33.8  5 0.5 1.77 1.85 0.415 0.73 41.7 0.77 38.9  60.5 1.71 1.81 0.412 0.70 45.9 0.75 43.9  7 0.5 1.50 1.75 0.412 0.62 49.50.72 48.7  8 0.5 1.34 1.61 0.417 0.56 52.8 0.67 53.2  9 0.5 1.44 1.690.410 0.59 56.2 0.69 57.8 10 0.5 1.06 1.28 0.415 0.44 58.8 0.53 61.3 110.5 0.90 1.18 0.409 0.37 61.0 0.48 64.5 Wash 1 0.00 0.00 0.390 0.00 61.00.00 64.6 12 0.5 1.03 1.040 0.403 0.42 63.4 0.42 67.3 13 0.5 0.91 0.9600.415 0.38 65.6 0.40 70.0 14 0.5 0.86 0.93 0.410 0.35 67.7 0.38 72.5 150.5 0.85 0.89 0.412 0.35 69.8 0.37 75.0 16 0.5 0.77 0.75 0.420 0.32 71.70.32 77.1 17 0.5 0.63 0.64 0.415 0.26 73.2 0.27 78.9 18 0.5 0.65 71.0%0.402 0.26 74.7 0.29 80.8 19 0.5 0.58 0.65 0.415 0.24 76.1 0.27 82.6 200.5 0.52 0.6 0.412 0.21 77.4 0.25 84.2 21 0.5 0.53 0.58 0.415 0.22 78.70.24 85.8 22 0.5 0.48 0.53 0.417 0.20 79.9 0.22 87.3 23 0.5 0.41 0.480.413 0.17 80.8 0.20 88.6 Wash 2 0.002 0.0024 0.540 0.00 80.9 0.00 88.624 0.5 0.37 0.43 0.410 0.15 81.7 0.18 89.8 Wash 3 0.0023 0.0019 0.5900.00 81.7 0.00 89.8 16% Solids on a weight/weight Basis gl Fe Ti TimeVolume Cumulative Cumulative Leach Hours Fe Ti Liters gm Extraction gmExtraction 0.25 7.07 4.00 0.010 0.07 0.2 0.04 0.1  1 0.50 8.16 4.520.380 3.10 9.3 1.72 5.9 0.25 4.77 4.63 0.010 0.05 9.4 0.05 6.0  2 0.505.78 5.44 0.402 2.32 16.3 2.19 13.3  3 0.50 4.74 4.54 0.418 1.98 22.11.90 19.6  4 0.50 4.16 4.13 0.419 1.74 27.2 1.73 25.4  5 0.50 3.62 3.630.418 1.51 31.6 1.52 30.5  6 0.50 3.30 3.27 0.425 1.40 35.7 1.39 35.1Wash 1 0.05 0.04 0.511 0.03 35.8 0.02 35.2  7 0.50 2.54 2.65 0.408 1.0438.8 1.08 38.8  8 0.50 2.08 2.41 0.417 0.87 41.4 1.00 42.1 Wash 2 0.500.08 0.09 0.450 0.04 41.5 0.04 42.3

EXAMPLE 3 Effect of Alkali Metal Halide

[0137] Experiments were conducted to evaluate the effect of an alkalimetal halide on the recovery of iron from ore using the sulfuric acidprocess of the subject invention. The results are shown in FIG. 4. Inthis case, the salt which was used was NaCl at 0%, 5%, 15% and 25%(w/w). These tests were performed using 200 gram per liter sulfuric acidsolution and no activated carbon at 100° C. on unground ore. Theaddition of salt speeds the reaction rate. However, at around 15 to 20%salt (150 to 200 grams per liter), NaCl appears to becomecounterproductive. The total percentage of iron leached actually fallsafter 15% NaCl is reached.

[0138] In a separate but analogous experiment to further demonstrate theeffect of an alkali metal halide on the leaching of iron out ofilmenite, the samples of 100 grams of ilmenite feed were leached with200 grams of sulfuric acid, 700 grams of water at 100° C. and varyingamounts of alkali metal halide for 72 hours. The amounts of alkali metalhalide were 0, 50, 150 and 250 grams representing 0, 5, 15, and 25%(w/w) alkali metal halide solutions. The leach rate of the iron wasobserved. The results are shown in Table 3. TABLE 3 The effect of alkalimetal halide on the sulphuric acid leaching process Fe Extraction TimeHours Fe mg/L Volume liters gm Cumulative No Alkali Metal Halide  0 860.700 0.06 0.2%  3 990 0.700 0.69 2.0%  6 1460 0.700 1.02 3.0% 12 30000.700 2.10 6.2% 24 4000 0.700 2.80 8.2% 36 6100 0.700 4.27 12.6% 4810200 0.700 7.14 21.0% 60 14600 0.700 10.22 30.1% 72 20700 0.720 14.9046.5% Wash 1 860 1.020 0.88 Wash 2 28 0.990 0.03 5% Alkali Metal Halide 0 100 0.700 0.07 0.2%  3 1110 0.700 0.78 2.4%  6 1860 0.700 1.30 4.0%12 3500 0.700 2.45 7.6% 24 4700 0.700 3.29 10.2% 36 6500 0.700 4.5514.1% 48 10100 0.700 7.07 21.8% 60 18600 0.700 13.02 40.2% 72 217000.790 17.14 56.5% Wash 1 1100 1.000 1.10 Wash 2 34 1.000 0.03 15% AlkaliMetal Halide  0 189 0.700 0.13 0.4%  3 1910 0.700 1.34 4.3%  6 25000.700 1.75 5.7% 12 4400 0.700 3.08 10.0% 24 5600 0.700 3.92 12.8% 367900 0.700 5.53 18.0% 48 12300 0.700 8.61 28 0% 60 18200 0.700 12.7441.4% 72 25300 0.720 18.22 63.9% Wash 1 1350 1.020 1.38 Wash 2 61 0.9900.06 25% Alkali Metal Halide  0 250 0.700 0.18 0.5%  3 2600 0.700 1.825.2%  6 4200 0.700 2.94 8.4% 12 9800 0.700 6 86 19.7% 24 11700 0.7008.19 23.5% 36 14500 0.700 10.15 29.1% 48 17300 0.700 12.11 34.7% 6018900 0.700 13.23 37.9% 72 18700 0.720 13.46 47.9% Wash 1 2960 1.0203.02 Wash 2 230 0.990 0.23

EXAMPLE 4 Effect of Grinding of Ore

[0139] In accordance with the subject invention, grinding of the ore canbe used to increase the reaction rate of leaching iron from ilmenite.This is shown in FIG. 5 and Table 4. Both tests were performed using a100 gram quantity of ilmenite placed in one liter of 200 grams per litersulfuric acid and 150 grams per liter alkali metal halide solutionheated to 100° C. The experiments were conducted on two samples of thesame ilmenite feed. One experiment used course ilmenite (100% retainedon a 75 micron screen) and the other experiment used fine ilmenite (100%passing through a 75 micron screen). The slurry was vigorously stirredfor 72 hours and the iron concentration periodically monitored. Theground ore (finer particle sized samples) had faster early and lateleach kinetics than the unground ore (coarser particle sized sample).The kinetics of the ore during the 5 to 25 hour time period was similarin both cases. TABLE 4 The Results of Feed Particle Size on theSulphuric Acid Leaching Process Time Fe Volume Fe Extraction Hours mg/lliters gm Cumulative Coarse Ilmenite  0 189 0.700 0.13 0.4%  3 19100.700 1.34 4.3%  6 2500 0.700 1.75 5.7% 12 4400 0.700 3 08 10.0% 24 56000.700 3.92 12.8% 36 7900 0.700 5.53 18.0% 48 12300 0.700 8.61 28.0% 6018200 0.700 12.74 41.4% 72 25300 0.720 18.22 59.3% Wash 1 1350 1.0201.38 Wash 2 61 0.990 0.06 Fine Ilmenite  0 2300 0.700 1 61 4.8%  3 50000.700 3.50 10.5%  6 5100 0.700 3.57 10.7% 12 5400 0.700 3.78 11.3% 246600 0.700 4 55 13.6% 48 20000 0.700 14.00 42.0% 60 24000 0.700 16.8050.4% 72 29300 0.815 23.88 76.3% Wash 1 1550 0.995 1.54 Wash 2 44 1.0000.04

EXAMPLE 5 Addition of Carbon Source

[0140] The addition of a carbon source in the form of activated carbonor graphite speeds the kinetics of the leaching reaction. The ratios ofcarbon to ore tried were 1:2 and 1:1. The results of tests carried outat 100° C. with 120 grams per liter salt and 200 grams per litersulfuric acid are shown in FIG. 6. The most cost effective carbon to oreratio will depend on final leach conditions. A person skilled in theart, having the benefit of the current disclosure, can identify theoptimal carbon to ore ratio for a particular process.

[0141] In a separate but analogous experiment to demonstrate the effectof the addition of a carbon source to the leaching of iron from ilmeniteto leave a TiO₂ concentrate residue, a 150 gram per liter alkali metalhalide solution was used, as opposed to the 120 grams per liter saltused in the aforementioned experiment. The experiment used ilmenitehaving a particle size wherein 100% of particles passed through a 75micron screen. A varying amount of coconut shell activated carbon wasplaced in each container. The same carbon to sample ratios wereevaluated as in the aforementioned experiment. The amounts were 0, 50and 100 grams of carbon for carbon to sample ratios of 0, 1:2 and 1:1,respectively. The slurry was vigorously stirred for 72 hours and theiron concentration periodically monitored. The results are shown inTable 5. The leach with the 1:1 ratio of carbon to feed material hadslightly better kinetics than the other two conditions. TABLE 5 TheResults of adding Carbon during the Sulphuric Acid Leaching Process FeeExtraction Time (Hours) Fe mg/l Volume (liters) gm Cumulative No CarbonSource  0 3000 0.700 2.10 6.0%  3 5700 0.700 3.00 11.5%  6 5900 0.7004.13 11.9% 12 6200 0.700 4.34 12.5% 24 14500 0.700 10.15 29.1% 48 250000.700 17.50 50 3% 60 27000 0.700 18.90 54.3% 72 36500 0.740 27.01 83.6%Wash 1 2060 0.995 2.05 Wash 2 60 1.010 0.06 1:2 Carbon:Ilmenite  0 41000.700 2.87 8.4%  3 8200 0.700 5.74 16.7%  6 11100 0700 7.77 22.6% 1212300 0.700 8.61 25.1% 24 20200 0.700 14.14 41.2% 48 24000 0.700 16.8049.9% 60 27000 0.700 18.90 55.1% 72 39300 0.695 27.31 89.3% Wash 1 32400.980 3.18 Wash 2 150 1.000 0.15 1:1 Carbon:Ilmenite  0 4800 0.700 3.3610.0%  3 8000 0700 5.60 16.7%  6 10100 0.700 7.07 21.1% 12 14500 0.70010.15 30.3% 24 24000 0.700 16.80 50.2% 48 27000 0.700 16.90 56.4% 6031000 0.700 21.70 64.8% 72 45400 0.550 24.97 93.5% Wash 1 6040 0.9905.98 Wash 2 355 0.990 0.35

EXAMPLE 6 Leaching of Copper and Nickel from Laterite Ore with aSulfuric Acid-Halide-Carbon System

[0142] This ore has an assay head of 2.36% Co, 1.26% Ni, 11.00% Fe,10.80% Mn. A sample of 100 grams of ground, −200 mesh ore was firsttreated with 200 grams of NaCl dissolved in 650 grams of water. Thewater was evaporated on a hot plate. This procedure is a speeded upsimulation of spaying a heap of ore with a salt solution and letting itevaporate naturally. The ore-salt solids were then slurried in 200 gramsof sulfuric acid in 700 grams of water solution. The stirred slurry wasbrought to 100° C. on a stirring hot plate, and then 100 grams of +65mesh, coconut shell, activated carbon was added. The test was run for 48hours with aliquots of solution taken at 1, 4, 6, and 24 hours. Theresults are shown in Table 6. The extraction of cobalt was probablycomplete within the first hour. The cobalt was probably precipitated bythe ionic strength of the solution and not recovered until the washsolution dissolved it. After 120 hours of leaching the ore under thesame conditions except for omitting the NaCl the Co recovery was 63.9%and the Ni recovery was 58.2%. TABLE 6 Percent Extracted from LateriteOre Leach Time (hrs) Co Ni Fe Mn  1 81.5% 67.5% 18.8% 63.4%  2 81.5%78.2% 25.5% 71.0%  4 81.5% 85.5% 31.5% 71.0%  6 81.5% 84.8% 36.0% 72.6%24 85.0% 87.5% 54.8% 73.3% 48 81.5% 86.2% 67.5% 71.0% Wash 1 13.1% 12.9%18.5% 11.5% Wash 2 5.4% 0.6% 2.2% 0.4% Final Liquors: 100.0% 99.6% 88.2%82.9%

EXAMPLE 7 Effect of Alkali Metal Halide on the Leaching of Nickel andCobalt from Laterite 1

[0143] Experiments were conducted on two samples of 100 grams oflaterite-1 feed, comprising 1.0 percent nickel and 0.1 percent cobalt ofa particle size of approximately 80% passing 75 microns.

[0144] In the first experiment the 100 g sample was leached with 200grams of sulfuric acid, 800 grams of water and no alkali halide at 100°C.

[0145] In the second experiment the 100 g sample was leached with 200grams of sulfuric acid, 800 grams of water at 100° C., and 200 grams ofalkali metal halide (sodium chloride).

[0146] Each experiment was run for a total of 6 hours with solutionsampling being carried out at 0.25, 0.5, 1.0, 2.0, 4.0, and finally 6.0hours. The results are shown in Table 7 and FIGS. 7 and 8. The secondexperiment that utilized the halide showed significantly better resultsfor both nickel and cobalt, and particularly cobalt. TABLE 7 Leaching ofNickel and Cobalt from Laterite-1 Ni Co mg/l Volume CumulativeCumulative Time Hours Ni Co liters mg Extraction mg Extraction No AlkaliMetal Halide 0.25 240.0 20.0 0.010 2.40 21.3% 0.20 20.3% 0.5 380.0 30.00.010 3.80 33.7% 0.30 30.5% 1 550.0 39.0 0.010 5.50 48.8% 0.39 39.6% 2740.0 53.0 0.010 7.40 65.8% 0.53 53.8% 4 820.0 64.0 0.010 8.20 72.7%0.64 65.0% 6 hr PLS 1060.0 75.0 0.780 826.80 94.0% 58.50 78.2% Wash 1116.0 40.0 0.670 77.72 26.80 Wash 2 12.0 1.3 0.700 8.40 0.91 IncludedAlkali Metal Halide 0.25 741.7 100.0 0.010 7.42 69.4% 1.00 97.7% 0.5867.2 100.0 0.010 8.67 81.2% 1.00 97.7% 1 958.4 100.0 0.010 9.58 89.7%1.00 97.7% 2 1026.9 100.0 0.010 10.27 96.1% 1.00 97.7% 4 1049.7 100.00.010 10.50 98.3% 1.00 97.7% 6 hr PLS 1050.0 100.0 0.765 803.25 98.3%76.50 97.7% Wash 1 103.8 10.0 0.670 69.57 6.70 Wash 2 60.5 5.7 0.59035.68 3.36

EXAMPLE 8 Effect of Alkali Metal Halide on the Leaching of Nickel andCobalt from Laterite-2

[0147] Experiments were conducted on two samples of 100 grams oflaterite-2 comprising 1.1 percent nickel and 0.1 percent cobalt feed ofa particle size of approximately 80% passing 75 microns.

[0148] In the first experiment the 100 g sample was leached with 200grams of sulfuric acid, 800 grams of water and no alkali halide at 100°C.

[0149] In the second experiment the 100 g sample was leached with 200grams of sulfuric acid, 800 grams of water at 100° C., and 200 grams ofalkali metal halide (sodium chloride).

[0150] Each experiment was run for a total of 6 hours with solutionsampling being carried out at 0.25, 0.5, 1.0, 2.0, 4.0, and finally 6.0hours. The results are shown in Table 8 and FIGS. 9 and 10. The alkalimetal halide (sodium chloride) test showed significantly better resultsfor both nickel and cobalt, particularly with regard to the speed withwhich full (100%) dissolution is achieved. TABLE 8 Leaching of Nickeland Cobalt from Laterite-2 Ni Co mg/l Volume Cumulative Cumulative TimeHours Ni Co liters mg Extraction mg Extraction No Alkali Metal Halide0.25 738.9 45.0 0.010 7.39 60.0% 0.45 51.6% 0.5 915.3 54.0 0.010 9.1574.3% 0.54 61.9% 1 1036.6 65.0 0.010 10.37 84.2% 0.65 74.5% 2 1113.880.0 0.010 11.14 90.4% 0.80 91.7% 4 1146.9 80.0 0.010 11.47 93.1% 0.8091.7% 6 hr PLS 1180.0 85.0 0.725 855.50 95.8% 61.63 97.4% Wash 1 264.721.0 0.530 140.28 11.13 Wash 2 51.8 4.1 0.695 36.02 2.85 Included AlkaliMetal Halide 0.25 906.5 80.0 0.010 9.07 76.4% 0.80 88.9% 0.5 1054.1 90.00.010 10.54 88.8% 0.90 100.0% 1 1149.0 90.0 0.010 11.49 96.8% 0.90100.0% 2 1159.5 90.0 0.010 11.60 97.7% 0.90 100.0% 4 1170.0 90.0 0.01011.70 98.6% 0.90 100.0% 6 hr PLS 1170.0 90.0 0.750 877.50 98.6% 67.50100.0% Wash 1 177.1 13.0 0.800 141.67 10.40 Wash 2 25.3 2.1 0.795 9.071.67

[0151] The above two experiments demonstrate the results for leachingboth the nickel and cobalt from two different nickel-cobalt lateritesamples.

[0152] The following procedure steps have been applied separately toeach of the laterite samples:

[0153] Step 1. A condenser is placed on the Erlenmeyer containing theslurry comprising the prescribed solution and laterite feed;

[0154] Step 2. The slurry is stirred vigorously with a magnetic stirrerfor the duration of the test with the temperature maintained at 100° C.;

[0155] Step 3. The test is sampled at predetermined times, eg., 15minutes, 30 minutes, etc., by pipetting 10 ml of the hot slurry from theErlenmeyer into a centrifuge tube and centrifuge at 4,000 rpm for 5minutes;

[0156] Step 4. The centrifuged timed leach solution is transferred intoa sample tube for later analysis;

[0157] Step 5. 10 ml of make-up leach solution is used to wash thecentrifuged residue back into the Erlenmeyer, while the Erlenmeyercontinues to be agitated at 100° C. on the hot plate;

[0158] Step 6. At the end of the test (e.g., 6 hours) the contents ofthe Erlenmeyer is poured into two centrifuge tubes, using an additionalvery small amount of distilled water to wash out any residue remainingon the inside lip of the Erlenmeyer, and then centrifuged;

[0159] Step 7. The centrifuged liquid contents (pregnant leachsolution—PLS) from both centrifuge tubes is decanted into a graduatedcylinder and allow to cool;

[0160] Step 8. Then having read the volume of PLS solution,approximately 20 ml is transferred into a sample tube and analysed fornickel and cobalt;

[0161] Step 9. Calculations are conducted to determine nickel and cobaltcontents of the liquors and comparisons made with the respectiveelemental assay values of the laterite ore feed;

[0162] Step 10. Nickel and cobalt can be totally or partially removedfrom the leach liquor by the method that makes the most economic sensefor any given plant. The methodologies available variously include, butare not limited to, precipitation of metallic salts by seeding, pHadjustment, or crystallisation; solvent extraction and electrowinning ofelemental metal; and ion exchange.

EXAMPLE 9 Leaching of Silver

[0163] This example shows the leaching of silver from a copper refinerypilot plant's slimes. The test was conducted at 100° C. with 200 gramper liter sulfuric acid and 200 gram per liter NaCl. Samples of 50 gramsof slimes were leached in 500 milliliters of solution. The leaching wasconducted for 48 hours. The results are shown in the Table 9. TABLE 9Results of 48 hour Leaching of Silver from Refinery Slimes Sample HeadAg, ppm Liquor Ag, ppm % Recovery Slime 1 14.75 18 80 Slime 2 58.9 59 65

EXAMPLE 10 Leaching of Molybdenum

[0164] A sample of molybdenum oxide ore with a head grade of 0.070% Mowas ground to minus 200 mesh and leached with agitation for 48 hours atroom temperature with a solution of 100 grams per liter sulfuric acidand 100 grams per liter sodium chloride. This leach recovered 89% of themolybdenum in the sample.

[0165] Another sample of unground ore from the same mine ore wasscreened to select the minus 18,850 plus 833 micron (minus ¾ inch plus20 mesh) fraction. This fraction was placed in a column and the same 100g/l sulfuric acid, 100 g/l sodium chloride was applied to the ore for 56days at 0.05 gallons per minute per square foot The leach solution wasrecirculated continuously. This leach scheme obtained 82% recovery ofthe molybdenum.

EXAMPLE 11 Two Stage Leaching of Different Metals into Two SeparateLeach Liquors

[0166] An oxide copper ore sample, ground to minus 200 mesh, with a headgrade of 0.91% Cu, 2.0 grams/ton Au, and 2.4 grams/ton Ag was leachedwith 100 gram per liter sulfuric acid for 72 hours in a stirred vesselat room temperature. The solid residue was then filtered and put intoanother vessel and leached for 30 hours at room temperature with asolution of 50 gram per liter of potassium bromide and enough sulfuricacid (6 ml) to adjust the pH to 1.0 with agitation. The results areshown in Table 10. TABLE 10 Percent Recovery in Stage Leach Stage Cu AuAg 100 g/l H₂SO₄ 71  0  0 KBr—H₂SO₄ 28 100 100

[0167] A person skilled in the art, having the benefit of the teachingsof this disclosure, can adjust the acid concentration and temperature toachieve complete recovery of the copper in the first stage whilemaintaining excellent recovery of the Au and Ag in the second stage. Thesubject method can also be readily adapted to heap leaching.

[0168] It should be understood that the examples and embodimentsdescribed herein are for illustrative purposes only and that variousmodifications or changes in light thereof will be suggested to personsskilled in the art and are to be included within the spirit and purviewof this application and the scope of the appended claims.

What is claimed is:
 1. A multi-stage method for recovering titanium andiron from a titanium and iron-bearing mineral feed, said methodcomprising: (a) contacting said feed material with an acid solution fora period of time sufficient to solubilize said titanium but insufficientto allow said titanium to appreciably precipitate, wherein a pulp and aleach liquor are formed; (b) separating said pulp from said leachliquor; (c) contacting said pulp with fresh leach liquor (d) repeatingsteps (a) and (b) until all economically feasible titanium is leached;and (e) recovering, separately, said titanium and said iron from saidleach liquors, wherein said solubilization step is carried out at atemperature of about 90° C. to about 105° C.
 2. The method according toclaim 1, wherein step (a) involves solids of between about 1 percent andabout 60 percent on a weight/weight basis.
 3. The method according toclaim 2, wherein said solids are of a particulate size such that saidsolids are capable of passing through a 75 micron sieve.
 4. The methodaccording to claim 1, wherein said acid is a halide acid.
 5. The methodaccording to claim 4, wherein said halide acid is selected from thegroup consisting of hydrochloric acid and hydrobromic acid.
 6. Themethod according to claim 1, wherein said acid is sulfuric acid.
 7. Themethod according to claim 1, wherein step (a) is carried out in thepresence of an alkali metal halide.
 8. The method according to claim 7,wherein said alkali metal halide is selected from the group consistingof NaCl, KCl, NaBr and KBr.
 9. The method according to claim 8, whereinsaid alkali metal halide is NaCl.
 10. The method according to claim 1,wherein said solubilization step is carried out at a temperature ofabout 40° C. to about 110° C.
 11. The method according to claim 1,wherein said solubilization step is carried out at a temperature ofabout 90° C. to about 105° C.